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By Lyall Zickrick

As a result of the studies on recrystallization and crystal growth made in this laboratory, certain theories have been developed. These are expressed briefly in a paper by Dean and Hudson.1 One of the principal points brought out in that paper is that in the deformation of a crystal, the energy supplied is distributed to the atoms of the lattice, probably by the forced formation of molecules. It is found that a Maxwellian distribution of the energy among these atoms will account for the grain growth at various temperatures. According to such a distribution the number of atoms with energy above a certain critical value, E, necessary for recrystallization at the absolute temperature T is given by n Ne(-E/kT) where N is the total number of atoms. It is apparent that if this line of reasoning is correct, a metal which has been cold-worked and reheated to a temperature below that of complete annealing will have the atoms with the higher energy recrystallized, that is, reduced to normal energy and there will remain only atoms below a certain energy. If this metal is again cold-worked the peak in the number-energy distribution curve will be displaced toward the regions of lower energies, that is, higher temperature of recrystallization. By a repetition of this process the position of this maximum and hence the recrystallization temperature of the metal could be raised considerably. In order to' check this theory, experiments have been made on copper wire. The wire was drawn from 1/4-in. rod to 22 B & S gage and then sample lots were annealed at 400 and 220° C., respectively. The wire annealed at 400° C. gave an elongation of 34 per cent., whereas that annealed at 220° C. gave only 23 per cent. These wires were then drawn by identical practice to 34 B & S gage and the elongation determined after an anneal of 3 mm. in an oil bath held at various temperatures.

Jan 1, 1927

By P. W. Andrews, R. B. Toombs

ECONOMIC STRUCTURE OF MODERN INDUSTRIAL ECONOMIES The role of minerals in modern industrial economies may be examined in several ways. There are relationships with the various sectors of the economy: the primary, secondary, and tertiary industrial groupings and activities. Mineral supply and demand activities can be measured in terms of national accounts criteria. The economic policy objectives of the modem industrial economy are based to a considerable extent on the needs of the nation for economic supplies of miner- al materials. There is an important role for minerals relative to national security in peace and in wartime and many national policies are directed toward ensuring adequate sup- plies of minerals. Environmental control in the modem industrial economy is increasingly concerned with matters related directly to the production and use of minerals and these matters, in turn, bear importantly on conservation in mineral supply and demand. Mineral polices include specific provision for research and development activities concerned with the production and the use of minerals. In all, national policy objectives will be shaped in considerable part by the circum- stances of mineral supply and demand: the relative degree of availability from domestic and foreign sources, the state of the mineral-processing industries, and the degree of development of the mineral-using industries, and per capita and total changes in requirements. Goods made from mineral raw materials are the basis of all advanced societies and although this is taken for granted, every country is continually dealing with problems of mineral supply and demand. Considering that in the first 50 years of this century the consumption of mineral materials in a highly Industrialized country such as the United States increased by six times while the population was doubling, it becomes apparent that if a country is to enjoy rapid economic growth it must have access to huge and ever-growing supplies of mineral materials at lowest possible cost. Adequate supplies of mineral materials and wise usage of mineral resources have always been among the most important determents of a nation's survival The importance of access to adequate mineral supplies is increasing every year and this has significant international, as well as national, implications because of the unique locational characteristics of mineral resources. It is generally acknowledged that no particular type of natural resource is essential to a high level of national income or to economic progress. A country's endowment of natural resources need not exercise a determining influence on the course of its national income over time, provided it is able to trade. What is required for economic progress is the availability of a substantial quanti-

Jan 1, 1976

[FUQUA, H. B. and Thompson, B. E.: Oil and Gas Development and Production in North Texas for 1933, 107, 320; for 1934, 114, 417 Oil and Gas Development in West Texas 114, 438 FURMAN, W. F.: Heat and Corrosion-resisting Alloys. Printed in Oil & Gas Jnl., June 13, 1935 Furnace-bottom sand, uniformity, 73, 390 Furnace control, Isley apparatus, M27, 502 Furnaces (See also Holt-Dern; Blast Furnaces; Roasters; Wirebar, etc.): annealing: electric, 89, 424 gas-fired, 89, 425 gas vs. wood for nonferrous metals, 89, 422]

Jan 1, 1936

PRESIDENT WILLIAM L. SAUNDERS,1 NEW YORK, N. Y. PAST PRESIDENTS CHARLES F. RAND,1 NEW YORK, N. Y. BENJAMIN B. THAYER,3 NEW YORK, N. Y. FIRST VICE-PRESIDENT PHILIP N. MOORE,1 ST. Louis, Mo. TREASURER GEORGE C. STONE,1 NEW YORK, N. Y. SECRETARY EMERITUS ROSSITER W. RAYMOND NEW YORK, N. Y. SECRETARY BRADLEY STOUGHTON NEW YORK, N. Y.

Jan 8, 1915

By G. R. Speich

The precipitation of the Feab and Fe,Ti Laves phases (MgZn, type, C14) from Fe-Nb and Fe-Ti solid solutions, respectively, has been studied in the temperature range 500" to 800°C using hardness measurements, light and electron microscopy, and X-ray and electron diffraction. The Fe-Nb alloys contained 0.96, 1.67, and 3.80 pct Nb and the Fe-Ti alloys 3.5, 5.8, and 8.8 pct Ti. In both systems only the equilibrium Laves phase is formed, with nucleation occurring at grain boundaries, at dislocations, and in the matrix. No cellular precipitation or transition phases were observed. The precipitation of the Laves phase in a fine dispersion leads to significant hardening in both systems. The hardness of the aged alloys is proportional to the reciprocal of the square root of the interparticle spacing. The maximum solid solubility of titanium in a! iron is 9.0 pct Ti at 1280°C. THE precipitation of the FezNb and FezTi Laves phases (MgZn2 type, C 14) from Fe-Nb and Fe-Ti solid solutions, respectively, has been investigated as part of a continuing study of the mechanisms of precipitation from ferritic binary iron-base alloys. The iron-rich portions of the Fe-Nb and Fe-Ti phase diagrams are shown in Fig. l(a) and l(b). The Fe-Nb diagram is essentially that given by ansen' with the more recent work of Hume-Rothery and Gibson2 added. Goldschmidt3 indicates that the FezNb Laves phase has a solubility extending from 32 to 55 wt pct Nb. The Fe-Ti diagram is also essentially that reported by Hansen.l The recent results of Nishimura and kamei, Murakami, Kimura, and Nishimura,' indicate that the FezTi Laves phase has a solubility extending from 22.5 to 30.5 wt pct Ti. The L/6 + FezTi eutectic temperature of 1298°C indicated by Kornilov and oriskina' has been adopted. The addition of niobium or titanium to iron lowers the 6 — y transformation temperature and raises the y — a! transformation temperature. This results in a closed y field in the Fe-Ti system, but the lower solubility of niobium in a! (6) iron results in a eutec-toid, 6 — y + FezNb, at 1200°C, and a peritectoid, y + FezNb —-a, at 989°C in the Fe-Nb system. The maximum solubility of niobium is reported to be 4.5 wt pct in 6 iron,' 1.0 wt pct in y iron,7 and 1.8 wt pct in a iron.' Some results from the present work indicate that the maximum solubility in a! iron is actually less than 0.96 wt pct Nb. The maximum solubility of titanium in a! (6) iron is somewhat uncertain since the early work of To faute and Butting-haus indicated a value of 6.0 wt pct while later work of Kornilov and Boriskina6 yielded a value of 14.0 wt pct. The present work indicates that the maximum solubility is actually 9.0 wt pct. In both systems the phase in equilibrium with the iron-base solid solutions is a Laves phase (Mg znztype, C14), FezNb or FezTi. Both of these Laves phases have a wide range of Solubility. The precipitation-hardening behavior of Fe-Nb or Fe-Ti alloys has not been studied extensively. Genders and Harrison' reported that binary Fe-Nb alloys hardened upon aging. Peter and Fisher7 studied the mechanical properties of Fe-Nb and Fe-Ti alloys quenched from different phase fields.

Jan 1, 1962

THE 143d* meeting of the American Institute of Mining and Metallurgical Engineers was held in New York, Feb. 19-22, 1934. It consisted of the annual business meeting, 45 technical sessions at which 217 papers were presented (more than ever before), two sessions of the Board of Directors, two meetings of Section delegates, three formal lectures, luncheon Monday, Tuesday and Wednesday for members and guests, three group dinners and five group luncheons, eight committee meetings, the annual reception and dinner dance, an informal dance, a smoke-dinner and the annual meeting of the Woman's Auxiliary. Attendance showed an improvement over the previous year, as follows: Total, 1534 in 1934, and 1475 in 1933; dinner dance, 543 and 426; smoker-dinner, 425 and 394; informal dance, 295 and 250; women's registration, 160 and 133; luncheons, 1034 and 734. The Coal Division held a general session and three sessions on Coal classification. It also met jointly with the Committee on Ground Movement and Subsidence. At the general session, John B. Dilworth presented a report of the Committtee on Methods of Valuing Coal Proper- ties and R. H. Sweetser a report of the Committee on Evaluation of Coal for Blast-furnace Coke. C. B Huntress, Executive Secretary of the National Coal Association, was the principal speaker at a luncheon on Monday, at which Eli T. Conner, Chairman of the Division, presided H. N. Eavenson, newly elected President of the Institute, and Edward W Parker, Secretary-Treasurer, of the Anthracite Institute, also spoke. The Executive Committee of the Division met Tuesday afternoon. The Sectional Committee on Classification of Coal met on Thursday morning and the Technical Committees on Scientific and Use Classification of Coal on Thursday afternoon. The Committee on Mining Methods held two general sessions, one for presentation of a half-dozen interesting papers on strictly mining subjects, the other on Aerial Transportation. It also participated jointly in a session on ground movement and subsidence with the committee of that name and the Coal Division The Institute's Standing Committee on Aviation had its inception at the session on Aerial Transportation

Jan 1, 1935

By H. H. Lowry

SCIENCE attracts the attention and interest of an individual or an industry in general only in proportion to the apparent direct application to its immediate welfare or benefit. Engineering accomplishments, as applied science, therefore receive more recognition than the scientific discoveries on which they are based. Even self-evident truth needs endless repetition before recognition: engineering as applied science cannot advance beyond the science on which it is founded. Scientific research, and discovery, is sometimes forced by the limitations recognized by those active in its development for the use of society and industry and sometimes results from an apparently superficially illogical curiosity about natural phenomena. Research has received inadequate support except in industries born in the laboratory, the chemical and electrical industries, from which many lessons might well be drawn by their elder associates. The rapid growth of these industries has been largely due to close cooperation of scientists and engineers. Most rapid advance is made when the research worker and the engineer keep in close touch with each other's ideas and problems. Many of the developments of modern industry are based on dis-coveries buried for years in scientific journals because the "time was not ripe" at their discovery. The situation with regard to utilization of coal is quite the opposite. The fundamental facts in regard to the scientific aspects of the problem of developing the energy from coal are scarcely known. The engineer, without much scientific cooperation, has accom-plished much in increasing the number of kilowatt-hours per pound of coal from 0.31 in 1919 to 0.69 in 1934, in increasing the 1000 gross ton-miles of freight hauled from 6.1 to 8.3 per ton of coal in the same period, in fuel economy in production of coke, and in developing the production of motor fuel in Germany and England by hydrogenation of coal. This lack of scientific research on the utilization of our most important natural resource, coal, has been due undoubtedly in large part to the reluctance of investigators to study the interrelationships involved in, the complicated processes in which coal is used. These complexities are made

Jan 1, 1936

By P. R. Hines

Diphenyl guanidine is used as an accelerator in vulcanizing rubber. Other rubber accelerators are also flotation collectors, e.g., dithiocarbamate, thiazole, and the xanthates. Urea and its derivatives are good flotation collectors,' and guanidine is the nitrogen analog of urea. These two characteristics suggested testing diphenyl guanidine as a flotation collector. Diphenyl guanidine has been tried in flotation work previously, but references give no details.'.' Bunker Hill Co. of Kellogg, Idaho, supplied the ore sample used in these tests. The sample was typical of the ore milled in 1949 and contained 4.52 pct Pb and 1.08 pct Zn. A flotation test made by the author with potassium ethyl xanthate as a collector, the same flotation reagents, and the same grind employed in regular Bunker Hill mill practice was the standard for comparing results with the compounds and derivatives of guanidine. The compounds and derivatives of guanidine are only fair collectors of galena. However, if potassium ethyl xanthate were used in the galena float, its presence later in the sphalerite float would make it impossible to introduce another type of collector and be certain which collector contributed the result, so the compound or derivative of guanidine under test was used in both the galena and sphalerite floats in Table I. On both the galena and the sphalerite in the Bunker Hill ore, the depressing action of sodium cyanide and zinc sulfate is much greater with collectors of the guanidine type than with the xanthates. If the depressing agents are left out when a guanidine type of collector is used, but a pH of 8.6 is maintained with either lime or sodium carbonate, the depressing action is approximately the same as in the standard xanthate test. Consequently the depressing agents, sodium cyanide and zinc, were not used in the galena floats of tests 1226, 1228, and 1286, Table I, so the galena floats would be on a comparable basis with the standard xanthate test. As shown by the tests in Table I, the loss in the tail when diphenyl guanidine is used is 8.3 to 9.3 pct, as compared to 17.6 pct with potassium ethyl xanthate, or a recovery of 2.3 to 2.0 lb more zinc. Concentrate grade is 13.7 pct with potassium ethyl xanthate and increases to 22.9 and 31.8 pct with diphenyl guanidine, showing a much higher selectivity. Unless the Barrett's No. 4 coal tar creosote is considered to be a collector, no collector other than diphenyl guanidine is present in the tests, Nos. 1226, 1228, and 1286. Tests 1218, 1253, and 1263 in Table I1 were run to check the effect of the depressing agents used in flotation of the galena (sodium cyanide and zinc sulfate) upon subsequent flotation of sphalerite by diphenyl guanidine. Urea was substituted for potassium ethyl xanthate in tests 1253 and 1263 because it is a good collector for galena and a poor one for sphalerite, and the effect of the depressing agents is more marked with urea than with potassium ethyl xanthate. Test 1205 checks the effect of urea alone without any depressing agents when used for the flotation of galena. Table I1 shows that diphenyl guanidine as a collector for sphalerite is compatible with other flotation reagents used in floating galena. When the depressing agents sodium cyanide and zinc sulfate are used in the galena flotation step, the amount of lime in the sphalerite flotation must be increased (compare 1253 and 1263). Table I11 gives some of the compounds of guanidine which were tested and compared. Some of the compounds and derivatives of guanidine* are solu- * Diphenyl guanidine is in commercial production and is available on the market. ble in water and others are not. Those tested were fed dry. When an organic collector is promising, its compounds and derivatives should be tested and their comparative collecting power recorded. Eventually, in this way, it may be possible to discover some of the essential chemical characteristics of a good collector. For example, it is interesting to compare the diphenyl and dibutyl derivatives in Table IV with those of urea in Table 111, Example 3, U. S. Patent 2,664,198. All are good collectors. Table I shows some of the effects of lime on recovery of sphalerite by diphenyl guanidine. Table V

Jan 1, 1960

By Fred Porter

Pollution problems from smelters may be unique in industry since the pollutants cover the wide range of SOx, particulates, carcinogens (arsenic), plus a number of heavy metals, the health implications of which are not clearly understood. In addition to this wide range of pollutants, many of the plants are extremely old; and fugitive emissions may, in fact, be the most significant air pollution problem. The Environmental Protection Agency, both in an in-house and contract effort, has assessed the quantity of pollutants emitted and evaluated the health impact. This paper will review the status of EPA's regulatory efforts in both new and existing smelters in light of these pollutant assessment studies.

Jan 1, 1976

By C. G. Fink, G. L. Putnam

The dissolution of gold by cyanide solutions was studied by determining the time required for the solvents to dissolve gold leaf. Minute traces, even 0.5 ppm, of sulphide ion retard the dissolution of gold, and this behavior cannot be accounted for by the presently accepted hypotheses involving oxygen-depletion or thiocyanate formation. On the other hand, traces of the salts of lead, bismuth, mercury, and thallium considerably accelerate the dissolution. The beneficial effect of lead is dependent upon the pH and cyanide ion concentration of the solution. ALTHOUGH cyanide solvents are used very ex-x*. tensively for the recovery of gold from ores and concentrates, the normal rate of solution of the precious metal as determined by Barsky, Swainson, and Hedley,1 and others2 is less than about 3 mg per sq cm of exposed gold surface per hour, corresponding to a corrosion depth of approximately 1.5 microns (0.00006 in.) per hr. Although the slow rate may be increased to some extent in certain cases, as, for example, when the the gold is partially imbedded in iron pyrites,3 the leaching operation is generally the most time-consuming step of the cyanide process. Frequently it is necessary for the leaching solvents to remain in contact with the ores for 50 to 200 hr in order to dissolve all of the gold particles.' Not only does the slow solution rate cause some increase in the capital costs, but the cyanide consumption also may be increased because of vaporization of hydrocyanic acid and secondary reactions with minerals such as malachite (CuCO3 • Cu(OH)2), and pyrrhotite, which often are associated with gold. It is not surprising that numerous attempts have been made to accelerate the dissolution of gold in cyanide solutions. Experimental Gold Leaf Test: The gold leaf test was apparently originated by M. Faraday, who, by means of it, discovered that dilute cyanide solutions would dissolve gold readily in the presence of oxygen. The usefulness and reliability of the test have been discussed elsewhere.5,6 Essentially, the method consists of determining the time required for 5-ml portions of cyanide solvents to dissolve squares of 23 carat gold leaf, 0.5 cm on edge, when shaken in 10-ml test tubes with the solvents. The weight of gold leaf per test was 0.051 mg as determined by the direct weighing of a leaf 8.6x8.6 cm (weight = 15 mg) and measuring the area of the test sample (0.25 sq cm). Gold leaf is of remarkably uniform thickness.6 Since gold leaf, like native gold, contains small amounts of copper and silver, the results are comparable with those one might expect in cyaniding gold ores. As six or more tests can be made simul- taneously, the relative efficiencies of the solvents can be evaluated readily. All dilute solutions of lead, bismuth, thallium and mercury salts were prepared by the ordinary dilution method used by chemists. Our C.P. sodium cyanide and buffering reagents had no detectable amounts of either sulphide ion or heavy metals. Soluble Sulphides in the Cyanide Process: The action of soluble sulphides is of interest in the study of the dissolution of gold in cyanide solutions. That gold ores which contain compounds of arsenic and antimony often are not amenable to direct cyanide treatment is well known, and it said that arsenic and antimony are "cyanicides."7 It is commonly believed that part of the sulphur content of such ores is soluble in alkaline solutions to give sulphide ion, which reacts with the free cyanide content of the leaching solutions to form inert thio-cyanates or reacts with the dissolved oxygen to form sulphites or sulphates.8 A purpose of this paper is to demonstrate that sulphide ion may behave in a manner that cannot be explained by either the oxygen-depletion or thiocyanate formation theories and to prove directly that concentrations of sulphide ion which cannot be detected even by colorimetric methods may retard the dissolution of gold.

Jan 1, 1951

By C. G. Fink, G. L. Putnam

The dissolution of gold by cyanide solutions was studied by determining the time required for the solvents to dissolve gold leaf. Minute traces, even 0.5 ppm, of sulphide ion retard the dissolution of gold, and this behavior cannot be accounted for by the presently accepted hypotheses involving oxygen-depletion or thiocyanate formation. On the other hand, traces of the salts of lead, bismuth, mercury, and thallium considerably accelerate the dissolution. The beneficial effect of lead is dependent upon the pH and cyanide ion concentration of the solution. ALTHOUGH cyanide solvents are used very ex-x*. tensively for the recovery of gold from ores and concentrates, the normal rate of solution of the precious metal as determined by Barsky, Swainson, and Hedley,1 and others2 is less than about 3 mg per sq cm of exposed gold surface per hour, corresponding to a corrosion depth of approximately 1.5 microns (0.00006 in.) per hr. Although the slow rate may be increased to some extent in certain cases, as, for example, when the the gold is partially imbedded in iron pyrites,3 the leaching operation is generally the most time-consuming step of the cyanide process. Frequently it is necessary for the leaching solvents to remain in contact with the ores for 50 to 200 hr in order to dissolve all of the gold particles.' Not only does the slow solution rate cause some increase in the capital costs, but the cyanide consumption also may be increased because of vaporization of hydrocyanic acid and secondary reactions with minerals such as malachite (CuCO3 • Cu(OH)2), and pyrrhotite, which often are associated with gold. It is not surprising that numerous attempts have been made to accelerate the dissolution of gold in cyanide solutions. Experimental Gold Leaf Test: The gold leaf test was apparently originated by M. Faraday, who, by means of it, discovered that dilute cyanide solutions would dissolve gold readily in the presence of oxygen. The usefulness and reliability of the test have been discussed elsewhere.5,6 Essentially, the method consists of determining the time required for 5-ml portions of cyanide solvents to dissolve squares of 23 carat gold leaf, 0.5 cm on edge, when shaken in 10-ml test tubes with the solvents. The weight of gold leaf per test was 0.051 mg as determined by the direct weighing of a leaf 8.6x8.6 cm (weight = 15 mg) and measuring the area of the test sample (0.25 sq cm). Gold leaf is of remarkably uniform thickness.6 Since gold leaf, like native gold, contains small amounts of copper and silver, the results are comparable with those one might expect in cyaniding gold ores. As six or more tests can be made simul- taneously, the relative efficiencies of the solvents can be evaluated readily. All dilute solutions of lead, bismuth, thallium and mercury salts were prepared by the ordinary dilution method used by chemists. Our C.P. sodium cyanide and buffering reagents had no detectable amounts of either sulphide ion or heavy metals. Soluble Sulphides in the Cyanide Process: The action of soluble sulphides is of interest in the study of the dissolution of gold in cyanide solutions. That gold ores which contain compounds of arsenic and antimony often are not amenable to direct cyanide treatment is well known, and it said that arsenic and antimony are "cyanicides."7 It is commonly believed that part of the sulphur content of such ores is soluble in alkaline solutions to give sulphide ion, which reacts with the free cyanide content of the leaching solutions to form inert thio-cyanates or reacts with the dissolved oxygen to form sulphites or sulphates.8 A purpose of this paper is to demonstrate that sulphide ion may behave in a manner that cannot be explained by either the oxygen-depletion or thiocyanate formation theories and to prove directly that concentrations of sulphide ion which cannot be detected even by colorimetric methods may retard the dissolution of gold.

Jan 1, 1951

By Robert Y. Grant

This study outlines the history of metal mining in Japan and the characteristics of the industry as they existed from 1925 to 1945. Mining and milling operations are described together with the role of the Japanese Government. A description of post-hostilities conditions and the present state of the industry is included. MANY of Japan's larger and better known mines date back to the early centuries of the Christian Era. The first stimulus to mining in Japan came through the growth of Buddhism, A. D. 600 to 700, and the attendant need for copper and other metals for temple ornaments and statues. During the Feudal period, 1333 to 1568, mine operations were expanded to meet the demand for ornament metal and for financing of campaigns. After removal of the seat of government from Kyoto to Tokyo, about 1600, the Tokugawa Shogunate placed great emphasis on mining, and many new mines were opened, some of which are still operating. Toward the end of the Tokugawa period, mining had declined sharply from its earlier peak. The easily worked, high grade portions of the orebodies had been exhausted, and the mining techniques then available could not exploit the deeper, lower grade ores. Similarly, extraction of metal from the sulphide ores presented a major problem. With the Meiji Restoration in 1868 and the rapid opening of the country to western ideas came a rejuvenation of the mining industry. Mining, milling, and smelting processes were greatly improved, resulting in the extension of development work into hitherto inaccessible portions of orebodies. The demand for metal during World War I stimulated activity throughout the industry; copper mining, particularly, expanded to a point where Japan ranked second among world producers.* During the depression following the war, mining underwent a severe contraction in activity, recovering slowly during the late 1920's only to undergo a further setback in the worldwide depression which started in 1929. Mineral output declined in 1930, but the beginning of Japanese expansion, 1931 to 1932, brought about an increase in the output of mine products starting between 1932 to 1934. Production continued to rise, the peak output being attained 1939 to 1943 in virtually all commodities. The greatest production of iron ore, chromite, and manganese was not experienced until 1944 when blocked imports, coupled with strong demands for steel, provided the needed incentive. Through necessity, Japan produced much of her mineral needs from her own mines at high cost during World War 11. Taking the 1932 to 1936 period as an average, however, Japanese mines produced only 16 pct of the iron available to industry, 8 pct of the lead, 33 pct of the zinc, and 68 pct of the copper. Nearly all of Japan's output of gold, silver, copper, lead, and zinc between 1925 and 1945 came from 30 to 49 pct of the mines for which production data are available; for example, 66 of 174 gold-silver mines accounted for 98 pct of the output of this group. Further analysis of mine output indicates that a large part of Japan's metal output was derived from a few large mines; about 10 gold-silver, 15 copper, 3 lead-zinc, 5 manganese, and 2 iron mines account for most of the production. Many of these mines have been operated for 200 to 400 years. With the

Jan 1, 1952

By G. H. Clamer

Page Bronze Bearing Metals. By G. H. Clamer............... 162 Pennsylvania Railroad Anti-friction and Bell Metals. By F. M. Waring .. 166 The Tin-plate Industry. By D. M. BUck. Discussed by G. H. ClameR, J. W. RICHARDS............................ 168 The Aluminum Bronze Industry. By W. M. Corse............ 171 Bronzes, Bearing Metals, and Solders. By G. K. BURgess and R. W. Woodward. Discussed by G. H. Clamer, R. T. Roberts......... 175 Babbitts and Solder. By G. W. Thompson................ 184 The Cadmium Supply of the United States. By C. E. Siebenthal. Discussed by M. L. Lissberger, C. W. Hill, F. F. ColcoRd.......... 185 Solder, Its Use and Abuse. By M. L. LissbergeR............ 192 Bronze Bearing Metals G. H. ClameR,* Philadelphia, Pa.—Unfortunately, prior to the war no serious attention was given to the conservation of tin, notwithstanding that this country is practically dependent upon outside sources for its entire tin supply. Tin is of vital importance in many industries, but it is surprising how many and how excellent are its substitutes when we become acquainted with them. Tin has always been a relatively high-priced metal, and it is part of the human attitude to associate high prices with high standards; it is not until the price of a commodity becomes well-nigh prohibitive that we hunt for substitutes, because the idea of substitution seems always to involve an assumption that the substitute must necessarily be inferior. The history of the development of bearing bronzes is a striking example of this policy. In the early days, copper-tin alloys were almost universally used, the idea then being prevalent, which is still held by many, that a bearing to resist wear must be hard, and the harder the better. The favorite bronze bearing contained 90 per cent. copper and 10 per cent. tin; frequently, in service which was considered severe, even higher proportions of tin were used. Such hard alloys have great resistance to compression, but as a rule they had a very wide factor of safety in this

Jan 1, 1919

I AM sure you know that it is necessary to draw the gold into wire in order to make gold cloth, gold embroidery, or gold filigree work. Just as it is beaten to form leaf for ornamenting pictures, so also can gold be drawn easily. The same is done with silver and tin, and I believe it could also be done with iron, copper, and brass. Although brass is not so

Jan 1, 1942

By D. S. Choi, D. L. McCain

A method of estimating the chain pillar size required to support the wide working faces of longwall or shortwall coal mining is presented. The pillar size is determined through the use of coal strength, entry geometry, panel width, and overburden pressure. Design analysis is focused on a three-entry development system in the Pittsburgh Seam but can be used for other development plans and other seams by substitution of the proper geometric and strength data.

Jan 1, 1981

POSITIONS VACANT (Under this heading will be published notes sent to the Secretary of the Institute by members or other persons.) Metallurgist wanted for research in iron and steel. Man with experience in mills preferred. Salary $4,000. No. 13. Opening for a few technical graduates to act as instructors in mining to men at the mines under the supervision of the State mine inspector. Salary $1,800 per year. No. 14. ENGINEERS AVAILABLE (Under this heading will be published notes sent to the Secretary of the Institute by members or other persons.) Mining engineer and metallurgist open for engagement, as superintendent of mines or constructing engineer. Technically educated, with 14 years' experience in United States and Latin America in concentration, amalgamation, cyaniding, constructing. Just finished building large modern all-sliming cyanide mill in Central America. No. 18.

Jan 4, 1914

IN CHAPTER I it was shown that the approximate market value of the copper produced at the twelve big Porphyry mines from 1905 to 1931, inclusive, was $2,821,300,000. This represented the refined metal ready for shipment to the consumer. In addition, at least $50,000,000 worth of gold and silver had been produced, more than $30,000,000 coming from the Utah mine alone. Consequently, the revenues totaled about $2,871,300,000, as shown in Table 58. Society paid this sum for a given quantity of metal that it wanted for industrial use. Where did the money go? Who were the direct beneficiaries of the creation of this new wealth? These questions this chapter will answer in a general way. Because some necessary data are not available, an exact tabulation in detail can not be made, but Table 58 is believed to reflect with reasonable accuracy the disposition made of the aggregate income from the twelve operations. The word "operations" is used advisedly because the complexity of corporate relations precludes the possibility of picking out companies, as could be done if each enterprise were an independent entity such as Miami, for example. Some explanation of the table and some elaboration of the items under "Outgo" is desirable. In the first place, it is assumed that the sum of surplus current assets existing in various treasuries at the end of 1931 was substantially equivalent to the “new” money that actually was put into the enterprises as a group. Less than $20,000,000 in new capital was realized from securities sold by the Utah Copper Co., the remaining funds coming from earnings. In respect to some of the others, such as Chuquicamata, Inspiration, and Andes, the amount of new money was much greater. However, of the two items-total new capital invested and total current surpluses-the former

Jan 1, 1933

By J. E. Dutrizac, R. J. C. MacDonald, T. R. lngraham

When sintered disks of synthetic chalcopyrite (CuFeS2) were leached in acidified aqueous solutions of ferric sulfate, the following reaction stoichiometry was obtained: CuFeS2 + 2Fe2(SO4)3 = CuSO4 + 5FeSO4 + 2S Over the temperature range from 50º to 94ºC, the reaction displayed parabolic kinetics. The parabolic rate constant for the dissolution of copper is given by the equation: log.k(mg2/cm4-hr)= 11.850 - 3780/T The activation energy for the dissolution process is 17 ± 3 kcal per mole. The parabolic kinetics have been attributed to the progressive thickening of a sulfur film on the surface of the chalcopyrite. When the leaching solutions contain less than 0.01 molar Fe+3 , the Fe concentration influences the rate of leaching, probably through a mechanism involving the diffusion of ferric sulfate through the sulfur layer. At higher Fe+3 concentrations, the rate control in the leaching. reaction has been attributed to the diffusion of ferrous sulfate through the sulfur. The rate of reaction is insensitive to changes in acid concentration and in disk rotation speed. ThE reaction of acidic ferric sulfate solutions with various sulfide minerals is of practical interest for both bacterial and heap leaching. This leaching medium is generally used with low-grade ores that cannot be treated profitably by conventional means. In both bacterial leaching1-3 and heap leaching, the active agent for sulfide dissolution is ferric sulfate. Although the reactions of ferric sulfate with chalcocite, covellite, and bornite have been investigated,4*7 the kinetics of leaching chalcopyrite with ferric sulfate have not been thoroughly studied. This paper reports a study of that reaction. EXPERIMENTAL Reagent-grade sulfur was purified by the method of Bacon and FanelliB and then it was vacuum-distilled to remove any soluble magnesium salts that had been introduced during the purification procedure.9 From stoichiometric quantities of the purified sulfur and hydrogen-reduced electrolytic copper sheet (99.90 pct Cu), CuS was synthesized at 450°C in a vacuum-sealed, pyrex vessel. About 24 hr was required for the completion of the reaction. A similar procedure involving hydrogen-reduced iron wire (99.90 pct Fe) was used to synthesize FeS1.002. A 2-furnace arrangement was required. The iron was heated to 800°C while the sulfur was maintained at about 400°C. Although the reaction to consume the sulfur was rapid, the material required additional heating (1 week) in a sealed silica ampoule at 800°C before it was homogenized. X-ray powder diffraction analysis confirmed that the copper sulfide was covellite and that the iron sulfide was troilite. The composition of the iron mineral was confirmed by wet chemical analysis. The two sulfides were ground to minus 100 mesh, weighed in equimolar amounts, mixed thoroughly, and pressed into pellets at 80,000 psi. The pellets were vacuum-sealed in pyrex ampoules and then sintered for 3 days at 550°C after an initial heating at 450°C for a few hours. The pellets were then cooled, polished with 3/0 emery paper, rinsed in acetone, and stored. The material had the characteristic brassy color of chalcopyrite and was shown by X-ray diffraction to be CuFeS2. Microscopic examination of the polished surfaces revealed small inclusions of pyrite (approximately 0.5 vol pct) as the only impurity. The presence of small amounts of a second iron compound will not alter the amount of dissolved copper but might increase the amount of ferrous ion slightly. It was calculated that dissolution of all of the pyrite and 100 mg of Cu (a typical value) would change the expected ferrous concentration by only 4 pct. Microscopic examination of a pellet after leaching revealed that the pyrite was not preferentially solubilized; only those pyrite grains at the surface were attacked. Hence, the pyrite is unlikely to alter the rate of copper dissolution. The chalcopyrite disks were about 1.7 mm thick and 27 mm in diam. They were about 80 pct of theoretical density, and for this reason their true reaction area was somewhat larger than the 5.8 sq cm area presented by the polished face. The disks were cemented to lucite cylinders in such a way that only the polished face was exposed. The disks were then leached by methods previously described.6,7 RESULTS AND DISCUSSION Stoichiometry and Kinetics. The initial experiments were directed to the problem of resolving the stoichiometry of the leaching reaction. Disks of CuFeS2 were leached at 80°C for various periods of time in acidified ferric sulfate solutions that were protected from oxidation by a cover of flowing nitrogen. When the disks had been partly leached, they were removed, their soluble salts were washed out, and then they were treated with CS2 in a Soxhlet extraction apparatus. The ratio of elemental sulfur to dissolved copper thus obtained was approximately 2 to 1. After the extraction of elemental sulfur from the pellet, the residue consisted of unreacted chalcopyrite only. For runs in which an appreciable amount of copper was dissolved, the ratio of ferrous ion to cupric ion in the solution was

Jan 1, 1970

AS If see it, then, the engineering advances of which yob have studied here at Golden, the continued growth of. industry that you see about you here in Colorado, and the country-wide progress that I have not time to discuss-these illuminating facts should clear away any doubts as to your future. You have little license to ask, "Where is the call for engineering or what the need? Yankee fashion I might ask you by way of reply, "Have all the mines been opened? Have all the power-houses and transmission lines been built? Have the last improvements in. mill and smelter been in-stalled?" We know that the general answer is No, and the rest of the answer, which tells where are those mines and power plants and what are the new methods and processes in mining and metallurgy, must be sup-plied by you engineers of the future. And anyone who has even casually observed what recent years have brought forth in our profes- sion cannot but envy you your job. Our reconnaissance excur-sion into your future should not be much longer extended, but before turning back there are four landmarks that I would like to point out-not lofty peaks but useful tie points that are worthy of your attention as you traverse the territory of your future work. Two of these are in fact low spots or depressed areas which call for engineering activity; these are commonly called "waste" and "average." The other two are of larger significance and are rather extensive regions in the domain you are to enter and occupy. These larger tracts bear the suggestive place names "fact-finding" and "human engineering" and perhaps will prove to be the heights from which the broadest outlook is to be had in this "pilgrim's progress:" In the first place, waste deserves your attention, al-though I realize that it is possibly the most unpopular word anyone hailing from Washington can publicly utter. Its unpopularity naturally follows from its overuse in the somewhat strong charges made against the industrial and business world some years ago, when the doctrine of conservation was first preached in the land. But waste is a word to be used in an engineering sense as well as for political effect. Properly used, waste is not a word to shy at-rather it is a challenge to the engineer. For example, at present, the loss of 1 kilowatt-hour to every 5 sold is a clarion call to the electrical engineer. Waste is your target to shoot at, the obstacle for you to surmount, the giant for you to slay.

Jan 7, 1928

By D. F. Kaufman, H. R. Spedden, A. M. Gaudin

MANY studies of comminution have been made to ascertain the size distribution of the product and to evaluate the work of comminution in the light of the size distributions of the feed and product. Up to now, these studies have been essentially statistical in character, that is, a certain lot of feed was subjected to comminution in some specified way, and the aggregate product was fractionated into sizes, thereby losing all knowledge of individual relationship of feed to product pieces. Radioactive tracers offer a means to do something in this respect which could not be done before, namely, to follow the rupturing of some particular piece in its normal environment of other pieces. That is, it permits going beyond the usual statistical limitations of size distribution studies to what may be termed a personalized or individualized study. The purpose of this paper is to present some preliminary experiments conducted with this tool. The method employed was to mark radioactively some constituent of a feed. It is possible, of course, to consider the preparation of two lots of material of which one is radioactive and the other is not, and to blend the two ahead of the comminuting step; but to do so is open to the objection that the two preparations may not be identical. Therefore a technique has been chosen that removes this objection by merely taking out a size fraction of a comminution feed, rendering that fraction radioactive by exposure to a neutron flux, and then by returning it to Table I. Size Distribution of Offspring Albite Particles Originally 28/35 Mesh and in Admixture with Other Sizes After Grinding 2 min in a Steel Ball Mill Specific Activity ' Cumu- Corrected Distrl- latlve Size for Back- butlon In Distri- Fractlon ground, Weight, Product, button, of Product, cpm/gm g Pctb Pct Mesh (A). (W) (P) (ZP) + 28 0 56.0 0 100.1 28/35 62.6 54.0 24.8 75.3 35/48 62.8 59.4 27.7 47.6 48/65 41.1 53.0 16.2 31.4 65/100 29.6 45.7 10.2 21.2 100/150 23.7 37.0 6.6 14.6 150/200 23.3 25.1 4.4 10.2 200/270 20.1 19.0 2.9 7.3 270/400 17.8 21.2 2.9 4.4 -400 22.9 25.2 4.4 — 100.1 a These activity determinations were made in rapid succession in the order given. The specific activity (Ao) of the active 28/35 mesh fraction of the feed was measured at the beginning, after the measurement on the 65/100 mesh size fraction of the product, and; The end. The decay-corrected activities at those times were 246.7, 241.0. and 236.9 cpm per gm. The weight (W0) of the active 28/35 mesh fraction in the feed was 55.0. b Example of calculation for P in the 65/100 mesh oroduct frac- A W tion; A = 29.6, W = 45.7, Ao = 242.7, Wo = 55.0: P = — x — Ao Wo = 0.102 = 10.2 pet. the remainder of the charge for the comminution experiment. A relatively simple procedure was developed by which albite, containing sodium, was activated in the M.I.T. cyclotron. The cyclotron makes highspeed deuterons which impinge on a beryllium target, thereby producing a concentrated neutron flux. The mineral was exposed to this flux for 2 hr. This treatment changed enough of the sodium to sodium 24 (14.8 hr half-life, 1.4 mev ß) as to make detection and measurement easy. The nuclear reactions taking place were: 11Na23 (n,?) 11Na24 (irradiation) 11Na24 ß,?,? 12Mg24 (decay) The detailed technique of the experimentation was as follows: 40 kg of hand-sorted, lump albite were crushed to pass 10 mesh. After careful mixing of the lot, a screen analysis was made. The whole lot of material was fractionated on standard Tyler screens from 14 down to 200 mesh. Samples for experiments were compounded from these fractions in accordance with the screen analysis. When it was desired to make an experiment in which, for example, the 28/35 mesh size fraction was to be studied, the blend of size fractions was made as indicated above, except that the 28/35 mesh size fraction was added only after irradiation in the cyclotron. The blended charge containing the activated albite was ground for 2 min in a laboratory ball mill with a steel ball charge of controlled size distribution. The ground product was carefully sized on a set of Tyler screens in a Ro-tap. Each size was analyzed for radioactivity by the use of an end-window Geiger-Mueller counter and standard scaling circuit. This analysis was carried out in detail as follows: a 20-g sample was placed in a Petri dish, packed carefully to obtain reproducible geometric distribution with reference to the Geiger-Mueller tube, and the activity was counted for a 2-min period. Several determinations of the activity of the active size fraction in the feed were made at various times to establish the decay in activity with time. Linear interpolation was used to evaluate the activity that the active size fraction in the feed would have had at any given instant. The ratio of the observed activity in a size fraction of the product to the activity that the active size fraction in the feed would have had at the same time gives the fraction in the product size that came from the irradiated size in the feed. The general formula for finding the distribution, P, of a specific individual size fraction in the feed

Jan 1, 1952